RU2353679C2 - Metals extraction from sulfide materials - Google Patents

Abstract

FIELD: metallurgy.

SUBSTANCE: invention relates to extraction method of ferrous metal from the sulfide material. Method includes stages of preparation acid aqueous halogenide solution, allowing oxidation potential, enough for oxidation of sulfide material and dissolution in solution of valuable metal, addition of the material into acid aqueous halogenide solution for that sulphide material would oxidised and valuable metal would dissolved and separation of valuable metal from oxidised sulphide material. Furthermore method of contaminant removing from polluted by admixtures sulphide material includes following stages: mixing of material in water solution, in which polyvalent variety with relatively high oxidation rate contaminant is oxidised, making it soluble in solution, with receiving of containing contaminants refined material and they are restored till relatively lower oxidation rate polyvalent variety and removing of contaminant from solution with simultaneous regeneration of polyvalent varieties till its condition with relatively high oxidation rate.

EFFECT: process simplification of consumption reduction of energy.

22 cl, 9 dwg, 25 tbl, 12 ex

Description

FIELD OF THE INVENTION

The present invention relates to a method for recovering metals, in particular valuable metals, such as gold, from a sulfide material. The method can be applied both to uncontaminated impurities and to contaminated impurities sulfide materials, including those that have a relatively high carbon content (the so-called "two-material"), or do not contain carbon, or have a low carbon content (the so-called "one-material material"). When the term “relatively high carbon content” is used in this description, it refers to the carbon content in the sulfide material, which is usually about 2% by weight.

State of the art

Significant deposits and many sulfide materials are found all over the world, including economically significant metals to be recovered, in particular valuable metals such as gold and silver. For example, there are significant deposits and reserves of pyrite ores, including gold and / or silver and other valuable metals, such as platinum and platinum group metals.

Some of these deposits are contaminated with difficult to handle contaminants such as arsenic, antimony, bismuth, or other heavy metals. Ore processing can also be complicated if high levels of carbon are present, since carbon is associated with valuable metals such as gold and has a high degree of affinity for them.

Commercially useful oxidation options for sulfide materials used in practice include calcination, pressure oxidation (POx), and biooxidation (Biox). In the Pox and Biox processes, a sulfate medium is commonly used.

When sulphide ores are fired, significant problems arise due to the release of sulfur-based gases toxic to the environment (the so-called SO _x gases). When arsenic is present in the ore, toxic substances such as arsenic trioxide are formed. For these reasons, there is a tendency all over the world to abandon the burning of sulfide ores.

To avoid problems associated with firing, oxidation of sulfide materials under pressure is used, but high pressures (usually more than 30 bar) and relatively high temperatures (more than 200 ° C) are required. Pressure oxidation is usually carried out in a sulfate-based solution.

US Pat. No. 6,461,577 discloses a biooxidation process for treating arsenic sulfide materials, wherein the sulfide material is subjected to a two-stage Biox process for dissolving arsenic. The leaching process scheme is complex due to the use of bioleaching bacteria. In addition, biooxidation is known to be a slow process.

US Pat. No. 4,553,305 discloses a leaching method for recovering copper and silver from sulfide ore using a combination comprising a solution of iron (II) chloride and compressed oxygen. While copper dissolves in liquor, silver does not leach out and remains in the solid residue from leaching. Then silver should be removed from the residue using sodium cyanide, which is an environmentally harmful leaching agent.

US Pat. No. 4,410,496 discloses a leaching method for recovering copper, lead, and zinc from a sulfide ore using a combination comprising a calcium or barium chloride solution and compressed oxygen. The precious metals contained in the ore again remain un leached and are present in the solid residue from leaching and should be removed separately.

US Pat. No. 4,658,829 discloses a leach method for recovering metals from sulfide ore containing arsenic and antimony. In this method, a bulk sulfide concentrate is obtained from an arsenic-containing sulfide ore. The concentrate is pulverized in excess calcium chloride solution. Immediately after receiving the concentrate, it is necessary to determine the total metal content and the composition of the concentrate. To prevent the formation of arsenic in the process of soluble compounds or toxic arsenic vapors, the concentrate is mixed with a pulp of a balancing solution containing a given concentration of copper, lead, zinc or a mixture thereof in the form of sulfides of these metals. In this case, the concentrate and pulp of the equilibrating solution are combined to form a reaction suspension having a given metal content such that the molar concentration of arsenic and antimony in the mixture is approximately equal to the molar concentration of copper, lead and zinc and is in the range of 60-40 or 40-60. Immediately after the mixture is properly balanced, it is heated and aerated under pressure to oxidize the metals into soluble components. In other words, balancing is important so that soluble arsenic compounds or toxic arsenic fumes are not formed.

Therefore, it would be desirable to develop a simple hydrometallurgical method for extracting a valuable metal, especially gold, from a sulfide material.

SUMMARY OF THE INVENTION

In accordance with a first aspect, the present invention is a method for recovering a valuable metal from a sulfide material, comprising the steps of:

preparing an acidic aqueous halide solution having an oxidizing potential sufficient to oxidize the sulfide material and dissolve the valuable metal in the solution;

adding material to an acidic aqueous halide solution to oxidize the sulfide material and dissolve the valuable metal;

separating valuable metal from oxidized sulfide material.

In the present invention, it was unexpectedly discovered that when sufficient oxidation potential is maintained in an acidic halide solution, the sulfide material can be oxidized simultaneously with the dissolution of the valuable metal in one step.

In addition, it was unexpectedly discovered that when sulfide material is contaminated with arsenic, antimony or similar impurities, a valuable metal can be dissolved, while arsenic and similar impurities can be simultaneously leached and precipitated in one stage without the need for a balancing stage of the previous or initial solution.

In this regard, in accordance with a second aspect, the present invention provides a method for recovering a valuable metal from an impurity-contaminated sulfide material, comprising the steps of:

preparing an acidic aqueous halide solution having an oxidizing potential sufficient to oxidize a sulfide material and dissolve a valuable metal in a solution, and having a pH at which arsenic precipitates;

adding material to an acidic aqueous halide solution to oxidize the sulfide material, dissolve the valuable metal, and precipitate arsenic;

separating valuable metal from oxidized sulfide material and precipitated arsenic.

The method of the first and second aspects also differs from the POx and Biox methods in that a halide-based leach solution is preferred than sulfate-based.

The authors also noted that halides (like cyanide) form strong complexes with valuable metals such as gold, and therefore can facilitate the dissolution of the valuable metal and subsequent extraction of the valuable metal, for example, adsorption on activated carbon. However, since halides are weaker ligands than cyanide, the authors have developed a method in which a sufficiently high oxidation potential (Eh) in an acidic environment (pH preferably <3) provides the ability to dissolve a valuable metal, similar to the solubility of cyanide.

Advantageously, the method can be implemented in a closed loop or in a recirculation mode with concomitant economic benefits (for example, simplicity, low energy consumption, maintaining material balance, etc.). The authors also observed that the method can be used to extract valuable metals from any sulfide material, including difficult to process ores and concentrates, such as “double-backed” materials having a relatively high carbon content (for example, carbon-containing arsenopyrites).

The solution containing the valuable metal is preferably separated from the oxidized sulfide material and arsenic is precipitated (if present) in the solid / liquid separation step and then the valuable metal is recovered from the solution in the extraction step of the metal, preferably by adsorption on activated carbon, usually in one or more columns containing activated carbon. After adsorption of the valuable metal onto activated carbon, the carbon is preferably removed and burned to recover the valuable metal or eluted with a cyanide solution and the eluate is fed to the electrolysis step to recover the valuable metal by electrodeposition. In this regard, the present method compares favorably with commercial methods used in practice, in which cyanidation of the oxidation residue is necessary to extract a valuable metal (gold), which requires a separate leaching scheme. In the present invention, the valuable metal is already dissolved in the liquor, therefore, cyanide leaching is not required. In addition, many reputable environmentalists are now demanding the destruction of residual cyanide, especially in areas of special ecological zones.

In the case of sulfide material not contaminated with impurities (for example, one-refractory pyrite ore, not contaminated with arsenic and other impurities), the sulfide material is usually oxidized at one stage. In the case of contaminated sulfide material (for example, one- or two-supported pyrite ore contaminated with arsenic and / or carbon and other impurities), the oxidation of sulfide material is usually carried out in two stages, although the dissolution of the valuable metal is achieved in the first of these stages.

Typically, the solution is recycled to the oxidation stage of the sulfide material after it has passed the metal extraction stage. The metal recovery step is preferably carried out “on line” after the solid / liquid separation step and before the solution is recycled to the step of oxidizing the sulfide material. The use of the term “on line” refers to a stage that is intended as part of a solution loop (ie, a “loop” ending in the recirculation of the solution to the oxidation stage of the sulfide material). In addition, metal extraction techniques other than adsorption on activated carbon, including ion exchange, solvent extraction, etc.) can be used.

In the case of a double-ore ore comprising carbon, it may be necessary to provide an additional separate metal recovery step (i.e., separate with respect to the recirculation circuit of the solution) designed to recover the valuable metal that comes with the solid particles of the material from the oxidation stage of the sulfide material. Such a separate step may be necessary due to the fact that a certain amount of a valuable metal (for example, gold) comes with carbon during the oxidation process and does not dissolve. In a separate metal recovery step, the traditional calcination or smelting process can be used, and after calcination, leaching (for example, using a solution from the oxidation stage of the sulfide material) may optionally be used to recover the valuable metal (e.g. gold) remaining in the calcined solid particles of the material.

Typically, the valuable metal to be recovered is gold, but it can be silver, platinum or another platinum group metal, the recovery of which usually justifies the economics of the process.

The aqueous halide solution is most preferably a soluble metal halide solution, typically having a halide concentration of about 8 mol / L. The halide is preferably chloride, but it may be bromide or a mixture of halides such as chloride and bromide.

The method is preferably carried out in such a way that the metal in the solution of the dissolved metal halide functions in the form of polyvalent species. In this case, polyvalent species are usually chosen that have both a relatively high oxidation state to participate in the oxidation of the sulfide material and a relatively lower oxidation state to which they are reduced during oxidation. Advantageously, the polyvalent species can then be regenerated to their relatively high oxidation state, after which the regenerated polyvalent species can be recycled to the oxidation stage of the sulfide material to participate in further oxidation. Advantageously, the regeneration of the polyvalent species occurs during the leach stage (s) so that the regenerated species can be recycled to the oxidation stage of the sulfide material as part of a preferred closed loop or in a process recycling mode with concomitant economic benefits (for example, maintaining material balance, simplicity low power consumption, etc.).

The metal in the metal halide solution is usually copper, but it can also be iron, etc. Any of these polyvalent species effectively acts as a means for electron transfer. For example, in a solution recycled to the stage of oxidation of a sulfide material, the metal is in its relatively high oxidation state (for example, Cu (II) or Fe (III) and after oxidation it is in its relatively lower oxidation state (for example, Cu (I ) or Fe (II)). At the leaching stages, the polyvalent species usually exist as a pair (that is, in their high and low oxidation states). However, other polyvalent species may be used, including possibly cobalt, manganese and vanadium.

When the sulfide material is arsenopyrite, when regulating the oxidation potential, arsenic can be leached into the solution in the first leaching stage. However, preferably, the pH of the solution is adjusted so that immediately after leaching, arsenic precipitates as iron (III) arsenate (scorodite). In addition, when the sulfide material is arsenopyrite, the pyrite component is preferably leached in a second leaching step, in which the pH of the solution is also adjusted to preserve the arsenic as a precipitate of iron (III) arsenate. Thus, arsenic is removed from the solid residue process in the solid-liquid separation step and does not interfere with the recovery of the valuable metal.

For a single-refractory pyrite material uncontaminated with impurities, the stage of oxidation of the sulfide material usually involves one-step leaching, in which the pyrite material is simultaneously oxidized and a valuable metal is dissolved.

Each leaching stage can be carried out directly or countercurrently and, therefore, one or more reactors can be used at each stage.

The entire solution from the first leaching stage is preferably fed to the second leaching stage.

When the sulfide material is contaminated, for example, with arsenic, in the first stage of leaching, the material is usually contacted with a solution at Eh sufficient to leach the contaminant and dissolve the valuable metal (e.g. gold), preferably at Eh of about 0.7-0.8 volts (relative to SHE). At the indicated Eh of the solution, the pyrite component of the material does not substantially leach. In the first leaching stage, the pH of the solution is preferably less than 1, but more than about 0.5, so that the contaminant is precipitated immediately after its leaching. In the first leaching step, the temperature of the solution is preferably about 80-105 ° C, more preferably 80-95 ° C.

In the case of sulphide material not contaminated with impurities (when one-step leaching is used) or in the second leaching stage used to leach the pyrite component of the sulphide material contaminated with impurities, the material is usually brought into contact with a solution having an Eh sufficient to leach pyrite, preferably Eh, about 0, 8-0.9 volts. The pH of the solution is usually also less than 1, but more than about 0.2 so that the contaminant is precipitated immediately after its leaching. For pyrite leaching, the temperature of the solution is usually the same or higher than the leaching temperature of arsenopyrite, usually it is from about 90 ° C to 105 ° C.

To achieve a higher Eh solution in one-step leaching or in the second leach stage, it may be necessary to add an additional oxidizing agent such as oxygen, air, chlorine gas, peroxide, hydrogen, etc. In order to achieve the optimum pH of the solution in order to maintain the contaminant in its precipitated form and to regenerate the copper (II) ion, it may be necessary to add an acid, such as sulfuric acid, and / or a base, such as calcium carbonate, to a single or second leaching stage for increase pH, otherwise arsenic and iron will not precipitate and to some extent will dissolve. In this case, either in the single or in the second leaching stage, the oxidation of the pyrite component of the material can produce a sufficient or excess amount of sulfuric acid. Alternatively, hydrochloric acid or any other acid that does not affect the chemistry of the process can be used.

While the solution separated after leaching is fed to recover a valuable metal, the separated solid residue is usually removed.

After the recovery of the valuable metal, the conditioning step of the solution is preferably used to remove (precipitate) the iron (III) sulfate and thereby control the level of such varieties in the process. Typically, at this stage, limestone and calcium carbonate are added to the solution to form a hematite / gypsum precipitate, which is then filtered and removed with a solid residue from the leaching step (s). However, the removal of iron (III) is preferably controlled by adjusting the addition of limestone to maintain a certain amount of iron in the solution, which, in turn, prevents the precipitation of copper (II) (that is, since iron precipitates at a lower pH than copper and buffers the pH, while precipitating, as a result of which it acts as a protection against copper deposition).

In the solid / liquid separation step, the solid residue is filtered from the solution, however, other separation techniques can be used, such as settling the solid / liquid mixture, evaporating the solution, centrifuging, etc.

When a high level of carbon is present in the sulfide material (e.g., 2-20 wt.% Carbon), a surfactant, such as a protective agent preventing the adsorption of valuable metals (such as gold) onto the solution during the oxidation of the sulfide material, can be advantageously carbon contained in the material. The protective agent is usually one or more organic solvents including kerosene, phenolic ethers, etc. Alternatively, activated carbon may be added to the solution to preferentially adsorb gold. The use of a protective agent or activated carbon can eliminate the need for a separate metal recovery step designed to separate the valuable metal that might otherwise enter the solid residue with carbon.

The most advantageous application of this method is the extraction of valuable metals from pyrite ores and concentrates, where the contaminant is usually arsenic, antimony, bismuth, mercury, cadmium, etc., which are naturally found in many deposits of pyrite materials.

Other economically significant metals can be extracted in the process, including copper, nickel, zinc, lead, etc. In addition, in some applications, a contaminant in itself may be desirable or necessary for recovery. For example, a contaminant can be economically valuable or harmful to the environment, which causes it to be removed from the contaminated sediment (for example, this can be the case for a contaminant such as antimony, bismuth, cadmium, etc.).

The method of the second aspect is used when the sulfide material is contaminated with arsenic, antimony or similar metals. In this method, the valuable metal is dissolved, while the contaminant is simultaneously leached and precipitated in one step, with no need for a step of balancing the previous or initial solution. In some applications, for example, when the contaminant needs to be extracted separately (for example, because it has some economic value) or simply as an alternative to the method of the second aspect, it may be desirable that the deposition of the contaminant occurs separately from the leaching of the contaminant.

Therefore, in accordance with a third aspect of the present invention, a method for removing contaminants from contaminated sulphide material is provided, comprising the steps of:

mixing the material in an aqueous solution in which multivalent species with a relatively high degree of oxidation oxidize the contaminant, making it soluble in the solution, and a refined material containing contaminants is obtained and reduced to a relatively lower oxidation state;

removing contaminants from the solution with the simultaneous regeneration of polyvalent species to their state with a relatively high degree of oxidation.

This method also makes it possible to advantageously recover metals, in particular valuable metals, such as gold, associated with contaminated sulphide material. In addition, by removing contaminants while regenerating polyvalent species, the method can be advantageously carried out in a closed loop or in a recirculation mode with associated economic benefits, such as simplicity, low energy consumption, maintaining material balance, etc.

In some other applications, for example, when a contaminant needs to be removed before conventional calcination or sulphide ore smelting, or simply as another alternative to the process of the second aspect, it may again be desirable for the deposition of the contaminant to occur separately from the leaching of the contaminant.

Therefore, in accordance with a fourth aspect, the present invention provides a method for removing contaminants from contaminated sulphide material, comprising the steps of:

mixing the material in an aqueous solution having an oxidizing potential, regulated essentially to oxidize only the contaminant to make it soluble in the solution, whereby a refined material containing contaminants is obtained;

separating the solution from the contaminant-containing refined material.

By controlling the oxidation potential, the method of the fourth aspect advantageously allows the contaminant to be kept in soluble form, which facilitates its subsequent removal (for example, in a separate precipitation step).

So, for example, when the material is arsenopyrite and the polluting impurity is arsenic, the oxidation potential can be regulated, for example, in the first leaching stage so that arsenic is oxidized and dissolved, and pyrite is not oxidized. In addition, in the method of the third and fourth aspects, immediately after dissolution and separation of arsenic, the remaining pyrite component can then be oxidized to a greater extent in the subsequent (eg, second) leaching stage.

The term “contaminant-containing refined material” includes a material in which the contaminant is not completely removed from it, but which has sufficiently low levels of contaminants so that it can be suitably further processed (for example, in kilns and smelters) or Comply with environmental standards for disposal. The method of the third and fourth aspects is typically used to treat pyrite ores or concentrates, where contaminants typically include arsenic, antimony, bismuth, mercury and cadmium. These contaminants naturally occur in many deposits of pyrite materials. The method of the third and fourth aspects can also be applied to the processing of difficult to process ores and concentrates, such as arsenopyrites, in particular double-refractory ores having a high carbon content.

In the method of the third and fourth aspects, the contaminant is usually removed from the solution by precipitation in a separate precipitation step by introducing an oxidizing agent into the solution. Advantageously, the oxidizing agent can simultaneously cause the oxidation of polyvalent species to their relatively high oxidation state. Then, after the deposition and removal of contaminants and the regeneration of polyvalent species to their higher oxidation state, the solution can be recycled to the leaching stage.

In the precipitation step, the pH of the solution is usually maintained at a value of about 1.5-3. The pH of the solution is usually maintained by adjusting the supply of the oxidizing agent and / or alkaline agent to the solution. When an alkaline agent is needed, an alkaline salt, such as calcium carbonate, calcium oxide, sodium carbonate, sodium bicarbonate, etc., is usually added.

At the deposition stage, the oxidizing agent usually causes the contaminant to precipitate by oxidizing it to a relatively less soluble form (for example, oxidizing arsenic from an oxidation state of +3 to +5) and at the same time oxidizing polyvalent species. The oxidizing agent may be air, oxygen, gaseous chlorine, hydrogen peroxide, etc. In pyrite ores, the contaminant is usually precipitated as an iron / contaminated oxidation product (for example, iron (III) arsenate when arsenic is the contaminant).

After sedimentation of the contaminant, Eh and the pH of the solution are usually restored to the levels necessary to leach the contaminant to allow the solution to be recirculated to the leach stage. This can be achieved, for example, by controlling the addition of an oxidizing agent after deposition of a contaminant.

In the method of the third and fourth aspects, the contaminant can be oxidized and leached into the solution in the process of single or multi-stage leaching. Typically, the leaching process includes a first leaching stage in which the oxidizing potential for oxidizing substantially only the contaminant is regulated, making it soluble in solution, and a second leaching stage in which the oxidizing potential for oxidizing sulfide in the contaminant-containing refined material is increased. In this case, it is preferable that the mass of the pollutant is oxidized and dissolves in the first leaching stage and the remaining contaminant can be oxidized in the second leach stage.

Each leaching stage can be carried out again directly or countercurrently and in this case one or more reactors can be used at each stage.

The contaminant-containing refined material is usually separated from the solution in the first leach stage and fed to the second leach stage. In addition, the solution is usually separated from the impurity-containing refined material after each leaching step to remove contaminants therefrom, usually by precipitation in the precipitation step.

When in the method of the third and fourth aspects, the sulfide material is a pyrite material (for example, arsenopyrite or other contaminated pyrites), in the first leaching step, the contaminant is usually oxidized in an acidic aqueous solution, which usually has a pH of less than 1.0 with an Eh solution sufficient to oxidation of contaminants in the solution, but not for leaching essentially pyrite, usually at Eh of about 0.7-0.8 volts (relative to SHE) and temperatures up to about 105 ° C. In the second leaching stage, the pyrite material is also oxidized in an acidic aqueous solution, which usually has a pH of less than 1.0, but with a higher Eh of a solution sufficient to leach pyrite, usually at Eh of about 0.8-0.9 volts and at temperatures up to about 105 ° C. In order to achieve a higher Eh of the solution in the second leaching step, an oxidizing agent such as oxygen, air, chlorine gas, hydrogen peroxide, etc. can be added to the solution. When necessary, an acid such as sulfuric acid may also be added.

In the second leaching stage, in order to maintain a low pH of the solution in order to oxidize pyrite and to dissolve the remaining arsenic, which is usually present in its oxidation state +5, it may also be necessary to add an acid such as sulfuric acid, hydrochloric acid or any other acid, which does not affect the chemistry of the process. However, the addition of acid may not always be necessary (for example, when sulfur present in the ore or concentrate is oxidized and therefore a sufficient amount of sulfuric acid is generated in the solution).

In the method of the first and second aspects, the solution recycled in the leaching and precipitation step is usually a solution of dissolved metal chloride having a chloride concentration of about 8 mol / l, and the metal in the solution of dissolved metal chloride functions as polyvalent species (as defined in the method of the first and second aspect).

When a high level of carbon is present in the sulfide material in the method of the first and second aspects (for example, 2-20 wt.% Carbon) to prevent adsorption of any dissolved metals (especially valuable metals such as gold) onto the carbon in the material during a step of oxidizing (leaching) the contaminant into the solution, a surfactant may advantageously be added as a protective agent. The use of a protective agent can eliminate the need for calcination used to separate the valuable metal from carbon.

Thus, in accordance with a fifth aspect, the present invention provides a method for treating contaminated sulphide material having a relatively high carbon content to enable extraction of a valuable metal contained in the material, comprising the steps of:

leaching the material in an aqueous solution, where the metal is leached into the solution, while the carbon in the material is masked to prevent the adsorption of valuable metal on it;

extraction of valuable metal from the solution.

The expression “relatively high carbon content” refers to the level of carbon present in the sulfide material, typically about 2-20 wt.%.

Carbon may be masked by the protective agent indicated above. The method of the fifth aspect may otherwise be as defined in the first to fourth aspects.

After sedimentation and removal of contaminants and regeneration of polyvalent species to their relatively high oxidation state, the solution is usually recycled to the leaching stage. Since polyvalent species are regenerated to their initial oxidation state (before leaching), they are ready to participate in further oxidation and leaching.

In the method of the third and fourth aspects, metal recovery steps may be provided for recovering metal leached into a solution with a contaminant and / or present therein from a refined material containing a residual contaminant.

For example, in the case of a double-bore ore comprising carbon, after the final leaching stage, a metal extraction step may be necessary to recover the metal present in a refined material containing residual impurity that is adsorbed, for example, onto carbon (e.g., valuable metal, such as gold, etc.). Due to the fact that in the case of a double-refractory ore, the contaminant is substantially removed from the sulfide material during leaching, the metal recovery step may include a conventional calcination or smelting process. After calcination, to recover the metal remaining in the calcined solid particles of the material (for example, where the metal is a valuable metal such as gold), chlorine leaching or cyanide leaching may optionally be used.

Alternatively or additionally, before (that is, between the oxidation and precipitation of the contaminant) or after (that is, between the precipitation of the contaminant and recirculation to the oxidation step) deposition of the contaminant to remove the metal leached into the solution in the leaching step, the metal recovery step may be necessary " on the line". The expression "on line" refers to the stage located in the recirculation circuit of the solution. On-line metal recovery typically involves the adsorption of metal in solution to coal in a column of carbon, usually with activated carbon. Alternatively, other metal recovery methods may be used, including ion exchange, solvent extraction, etc.

Typical metals recovered in the process of the third and fourth aspects include valuable metals such as gold, silver, platinum or other platinum group metals, the extraction of which usually justifies the economics of the process. However, alternative or additionally, other economically significant metals may be recovered, including copper, nickel, zinc, lead, etc. In addition, in some applications of the method of the third and fourth aspects, it may be desirable or necessary to extract the contaminant itself. For example, a polluting impurity can be economically significant or harmful to the environment, which leads to its extraction from the residue contaminated with impurities (for example, this may be the case for such a contaminant as antimony, bismuth, cadmium, etc.). When the contaminant constitutes the “metal” to be recovered, after the contaminant has been deposited, a step for recovering the contaminant can be additionally or alternatively provided.

Prior to metal recovery, in the method of the third and fourth aspects, a series of material separation steps are usually provided for separating the impurity-refined material from the solution. In such a case, after the first leaching step, the solution is usually subjected to a thickening step to thicken and separate the contaminant-containing refined material from the solution. Usually, after the second leaching stage, the contaminated refined material is filtered out of the solution, however, other separation methods can be used, such as settling the solid / liquid mixture, evaporating the solution, centrifuging, etc.

Thus, after each of the first and second stages of leaching, the separated solution is usually fed to remove contaminants, while the separated refined material, if necessary, can be submitted for metal recovery (for example, in the case of double-backed pyrite) or for removal.

In addition, in the method of the third and fourth aspects, after the contaminant deposition step, a contaminant separation step is typically provided for removing the contaminant from the solution before recirculating the solution to the leaching step (or before the on-line metal recovery step). In this case, after sedimentation of the contaminant, a solid / liquid separation step is usually used, which may be facilitated by filtration or other separation techniques.

Brief Description of the Drawings

Although any other forms may be included in the scope of the present invention, preferred forms of the invention will now be described only by way of examples and with reference to the accompanying drawings, in which

Figure 1 schematically depicts a General view of the methods of POx and Biox of the prior art in comparison with the preferred method (IRGP) in accordance with the present invention, which are designed to extract valuable metal from a sulfide material;

Figure 2 schematically depicts a General view of the process for the first method of extraction of valuable metal (gold) from contaminated with impurity sulfide material (arsenopyrite - FeAsS);

Figures 3 and 4 depict graphically the extraction of gold and iron and the time dependence of the Eh solution for the various stages of IRGP;

Figure 5 schematically depicts in general terms a process diagram for the second method, illustrating the removal of contaminants from a sulfide material and the recovery of a valuable metal from a sulfide material;

6 schematically depicts a process flow diagram for a preferred method for removing contaminants from a single-refractory sulfide material and recovering a valuable metal from a sulfide material;

7 schematically depicts a process flow diagram for a preferred method for removing contaminants from a bicomponent sulfide material and recovering a valuable metal from a sulfide material;

Fig. 8 is a graphical plot of various parameters of the leaching solution (arsenic) in the first stage versus time (reaction time) for the method of Figs. 6 and 7;

Fig.9 is a graphical dependence of various parameters of the leaching solution (pyrite) in the second stage from time to time (reaction time) for the method of Fig.6 and 7.

Embodiments of the invention

Before describing the preferred methods of the invention using details and examples, preferred methods in accordance with the present invention will first be described in general terms and in the context of the prior art and Pox and Biox methods and with reference to FIG.

Preferred methods in accordance with the present invention are indicated in this detailed description as "Intec Refractory Gold Extraction Method" (IRGP). These methods were developed as an alternative based on a halide for the extraction of gold from refractory sulfide mineral deposits. The known treatment of such deposits is usually carried out by flotation of the crushed ore to obtain a concentrate which is subsequently treated to oxidize sulfide minerals, preferably to sulfate, with the culmination of the treatment being the recovery of gold from the oxidation residue using cyanide.

Practical commercially available oxidation options for sulfide minerals include firing, oxidation under pressure (POx), and biooxidation (Biox). The essence of IRGP through comparison with current hydrometallurgical practice (POx and Biox) is schematically depicted in Figure 1. IRGP compares favorably with the hydrometallurgical variants of POx and Biox in that a halide medium is preferred over a sulfate medium. Gold is insoluble in sulfate, while halides, like cyanide, form strong complexes with gold, which facilitate its dissolution and subsequent extraction by adsorption onto activated carbon. Since halides are weaker ligands than cyanide, an acidic medium (pH <2) and higher solution temperature and potential (Eh) are used to achieve the same degree of gold recovery.

In the treatment of refractory sulfides, a halide medium with controlled oxidative potentials of the solution provides the possibility of oxidation of arsenic and sulfide and the dissolution of gold. After separation of the gold-loaded solution from the pulp of the oxidized sulfide mineral, the dissolved gold is able to be adsorbed onto activated carbon, which is subsequently burned or eluted with cyanide to finally recover the metallic gold by electrodeposition. In contrast to traditional commercial practice, IRGP does not require cyanidation of the oxidation residue to extract gold, which requires a separate, specially designed leach circuit and possibly also the additional cost of destroying the cyanide residue.

There are a number of factors that can make gold ore refractory, as shown in the following table:

Type of Reasons for persistent properties Release Physical locking in silicates, sulfides, carbon, etc. Occlusion Passivation due to the formation of a chemical layer Chemistry The formation of gold-containing compounds, for example, gold tellurides and aurostibite Substitution Elemental substitution of gold in a mineral lattice, for example, a “solid solution” of gold in pyrite ores Adsorption Adsorption of dissolved gold by "active" carbonaceous material in ore pulp

IRGP was developed specifically for the processing of concentrates obtained from refractory ores, which are included in the last two categories “substitution” and “adsorption”. The majority of the world's gold reserves fall into the two indicated categories, in which iron sulfides, such as arsenopyrite and pyrite, are prevalent, either individually or more frequently in combination. IRGP is also applicable when carbon is also present in the ore.

The IRGP process and chemistry are now described for processing refractory gold-bearing concentrates containing the following types of minerals:

1. Arsenopyrite.

2. Arsenopyrite + pyrite.

3. Arsenopyrite + pyrite + carbon.

Arsenopyrite Oxidation Chemistry

Arsenic in refractory gold-bearing concentrates is present mainly in the form of arsenopyrite (FeAsS). Gold is usually "locked up" in such arsenopyrite mainly in the form of varieties bound in the crystal lattice, which are often attributed to solid solution rather than natural gold. Therefore, the release of gold requires the complete destruction of the arsenopyrite lattice.

The destruction of the arsenopyrite lattice in IRGP was achieved by chemical oxidation in accordance with the following general reaction:

Oxygen did not oxidize arsenopyrite directly, but acted through several intermediate steps, as its solubility in process liquor was extremely low.

Oxygen was supplied directly from air sprayed into the liquor at atmospheric pressure, and was originally used to generate a soluble oxidizing agent in the form of a copper ion (Cu ²⁺ ) in accordance with the following reaction:

This reaction proceeded at the interface between air bubbles and technological liquor. Then the divalent copper ion oxidized arsenopyrite in accordance with the following reaction:

The reaction products, which are ferrous iron and monovalent copper, were then oxidized by additional spraying of air in accordance with reaction (2) and the following reaction:

In the presence of a ferric ion, arsenic acid readily formed an insoluble ferric arsenate in accordance with the following reaction:

Ferric arsenate was able to form in a strong chloride electrolyte and, under the operating conditions used in IRGP, was usually crystalline and stable in the environment, which made it easy to separate.

The action of the Cu ²⁺ / Cu ⁺ pair was supplemented by the Fe ³⁺ / Fe ²⁺ pair, since a small background concentration of iron was always present in the technological liquor. The potential reached under the influence of Cu ²⁺ and Fe ³⁺ was in the range of 850 mV (in comparison with SHE) in the presence of oxygen. This potential was sufficient for the dissolution of gold due to the stabilization of gold due to the formation of a chloride complex in accordance with the following reaction:

When bromide was present in the process liquor (for example, if it was added intentionally), a gold-bromide complex was also formed in accordance with the following reaction:

The oxidation was carried out at a temperature of 90-95 ° C in 8M chloride electrolyte containing 20-40 g / l of Cu ^{2+ ion} and 2-5 g / l of Fe ³⁺ ion.

Chemistry of Pyrite Oxidation

The oxidation of pyrite (FeS ₂ ) in IRGP was achieved through the same sequence of intermediate reactions that was used to oxidize arsenopyrite, in accordance with the following general reaction:

It was noted that all pyrite sulfur was completely oxidized to sulfate, in contrast to arsenopyrite sulfur, which was oxidized only to an elementary state.

Pyrite is more refractory than arsenopyrite, and therefore, finer grinding was used to achieve acceptable reaction kinetics, which are explained below. However, individual pyrite samples showed variable chemical activity, which suggested the effect of the substitution of part of the sulfur in the crystal lattice with arsenic. Such pyrite is often called arsenic pyrite, and the higher the arsenic contamination, the greater the chemical activity of pyrite approached that of actual arsenopyrite, and the As / S ratio is unity.

The reaction proceeded through the Cu ²⁺ / Cu ⁺ pair, as in arsenopyrite, at a temperature of 90-95 ° С in the same liquor that was used for the oxidation of arsenopyrite, in accordance with the following reaction:

Cu ⁺ and Fe ^{2+ were} oxidized with additionally sprayed oxygen in accordance with reactions (2) and (4). The ferric sulfate formed is precipitated as hematite and gypsum by the addition of limestone at a pH of about 1-1.5 in accordance with the following reaction:

The maintenance of the soluble iron concentration in the range of 2-5 g / l was regulated by the addition of limestone, which prevented the precipitation and loss of divalent copper with the residue from leaching.

The size of the grinding concentrate

Concentrates for use in IRGP are usually obtained with a grinding size of 80% concentrate in the range of less than 70-100 microns. Tests showed that the kinetics of the reaction increased significantly when the concentrates were regrind to a finer grinding (depending on the properties of each individual concentrate), and re-grinding was usually used in the first version of the method (disclosed below). When arsenopyrite was only a gold-bearing mineral, it was proved that a size of 80% of the concentrate, equal to less than 30-40 microns, was adequate to achieve high gold recovery and an acceptable leaching time.

When gold was "locked" in pyrite, the size of the grinding depended mainly on the chemical activity of pyrite, which, as explained above, changed significantly. For highly active pyrite, grinding was used, which was used for arsenopyrite, but for examples of more refractory pyrite, finer grinding was required. In the case of extremely refractory pyrite, ultrafine grinding was sometimes used, with 80% of the concentrate having a grinding size of less than 6-10 microns. The authors also note that the ultrafine grinding technology developed over the past 10 years until the moment when ultrafine grinding mills began to operate successfully in mines around the world.

Gold recovery

A gold-containing leach solution was passed through columns containing activated carbon onto which gold was adsorbed. The duration of gold adsorption was 10-15 minutes, which is typical for traditional practice using cyanide compounds. The load of gold on carbon was usually 2-5 wt./wt.%, Due to the relatively high concentrations of gold in solutions (usually 10-100 mg / l), which is a consequence of high-grade gold in concentrate. The extraction of gold under such a load was carried out due to the destruction of carbon by burning in a furnace. At a lower load, elution with cyanide followed by coal reactivation was more economical.

Cleanliness management

In addition to the main contaminants (such as arsenic, antimony, etc.), impurities (such as Cd, Mn, Mg, etc.) are present in the initial concentrate, which do not adversely affect either the leaching operation or the operation deposition. Nevertheless, to prevent their accumulation in the technological solution over time, the impurity control method is used. Prevention of their accumulation was achieved by precipitation from the drain of the regenerated solution of divalent copper with purified brine returned to the process. It is important that no wastewater forms in the IRGP and that all impurities are removed as solid by-products.

To bring the pH to 3.5, limestone was added to the drain, which precipitated residual iron and copper, which were removed by filtration and recycled to the liquor. Then impurities such as Cd, Mn, Mg were removed by adding slaked lime at pH 9.0, and insoluble oxides formed, which were removed by filtration and removed.

In the context of the process equipment, the IRGP was similar to the Biox methods in that atmospheric pressure was used, but the process time was much shorter and was usually 6-20 hours. During pyrite oxidation, the leaching temperature was higher than in the Biox method, and the elimination of the oxygen supply unit (which is used for POx) was achieved when the concentrate supplied to the process was finely divided, usually to an ultrafine size, for example, to a size of less than 10 microns . The materials from which the process equipment was made were fiber-reinforced plastic, rubber-lined steel and titanium.

Arsenopyrite plus pyrite plus carbon (double back)

The influence of carbon in the processing of gold-bearing concentrates largely depended on their grade and activity. When the carbon content was in a low range, either organic additives (protective agents) were used to suppress the adsorption of gold, or activated carbon was added to the liquor for preferred adsorption of gold (CIL - liquor in coal). Thus, in such cases, the oxidation of arsenopyrite was as described above.

However, when the initial carbon content exceeded 3-5%, the efficiency of suppressing the adsorption of gold or CIL was significantly reduced, as the so-called “leaching” of gold increased. In this case, the destruction of carbon by calcination was the main processing option practiced in the prior art methods in this field. Calcination is a relatively complex process, as the conditions of calcination affect the recovery of gold from the cinder. In addition, the optimal conditions for burning pyrite differ from those for arsenopyrite, which inevitably entails the use of a two-stage firing process.

The use of IRGP before firing provided selective leaching of arsenic and sulfur and simplified subsequent firing, which in this case became a simpler one-step process. In addition, the removal of arsenic and sulfur reduced the formation of exhaust gas from kilns, because As ₂ O ₃ and SO _{2 were} formed to a much lesser extent. The use of IRGP resulted in a significant reduction in capital and operating costs at the firing stage.

The first and second variants of the method

In the processing of refractory sulfides in the first embodiment of the method in accordance with the invention, the use of a halide medium with a certain oxidation potential of the solution made it possible to oxidize sulfides simultaneously with the dissolution of gold (the so-called all-in-one process). In the second variant of the method in accordance with the invention, the use of a halide medium with other solution parameters ensured the oxidation of the polluting impurity (for example, arsenic, etc.) before oxidation of the sulfide, usually with some dissolution of gold, while the separation of the polluting impurity and further extraction of gold were carried out separately and sequentially. Now, the first embodiment of the method in accordance with the invention will be described in detail, and then the second embodiment of the method in accordance with the invention will be described in detail.

The first variant of the method

When describing the first variant of the method, reference will be made to FIGS. 2-4 and Examples 1-3.

As follows from Figure 2, it schematically depicts a method for extracting gold 10 from single-refractory pyrite. Served in the process concentrate 12, containing valuable metals, obtained by mining, grinding and subsequent flotation of sulfide ore. The concentrate is usually a gold-containing arsenopyrite (when it has a high carbon content, it becomes bicarbonate). The concentrate is ground in a special ball mill, usually to an ultra-thin size of less than 10 microns. The ground concentrate is then fed to the first leaching stage in the form of the leaching stage of arsenopyrite 14.

During the leaching of arsenopyrite 14, an acidic medium is maintained (pH is preferably less than 1, since at low pH the leaching of arsenopyrite is enhanced). An acidic medium can only be achieved by recirculating the solution, or an acid-free acid (such as sulfuric or hydrochloric acid) can also be added. The Eh of the leach solution is usually maintained at a value of more than 0.4 volts to promote oxidation of the arsenopyrite component of the material and the dissolution of gold. The leaching temperature is maintained at a value of about 80-95 ° C.

The leached material is then fed to the second stage of leaching of pyrite 16, where an oxidizing agent (such as oxygen, air, chlorine, hydrogen peroxide, etc.) is added to increase the oxidation potential of the solution and, consequently, oxidize the pyrite. To maintain the arsenic in precipitated form in the second leaching step, it may be necessary to add an acid (such as sulfuric acid) or a base (such as calcium carbonate) to maintain the pH of the solution above about 0.2.

The process solution is usually an aqueous solution of divalent copper chloride having a chloride concentration of 8 mol / L. At both stages of leaching of arsenic and pyrite, the divalent copper ion oxidizes the sulfide material and is reduced to a monovalent copper ion (equations (2) and (8)). The divalent copper ion is also regenerated in an acidic oxidizing medium (equations (3) and (9)). Thus, in the method, copper acts as a means for electron transfer, existing in the form of a pair of Cu ²⁺ / Cu ⁺ . This function can be performed by other agents, including iron, cobalt, manganese, vanadium, etc.

When the sulfide material has a high carbon content (for example, up to 3-5 wt.%), A masking surfactant can be added to the solution at stages 14 and 16 to prevent the adsorption of gold (or other valuable metal) leached into the carbon onto the carbon. The surfactant is usually an organic protective agent, such as kerosene, phenol ether, etc. Alternatively, to preferentially adsorb gold and subsequently remove it, activated carbon can alternatively be added.

The authors found that in the first embodiment, at the stage of leaching of arsenopyrite 14 at a regulated pH of less than 1, but higher than the pH at which arsenic dissolves, at a regulated relatively moderate Eh of about 0.7-0.8 volts (compared with SHE) and at relatively at low temperatures (80-95 ° C), the material can be leached and gold dissolved, without oxidizing pyrite sulfide to sulfate.

The oxidation conditions used at the stage of leaching of pyrite 16 are more stringent than the conditions at the stage of leaching of arsenic 14. In this case, the oxidizing agent is sprayed into the solution so that the oxidation potential Eh rises to about 0.85 volts. In addition, in the second leaching stage, it may be necessary to increase the temperature of the solution to about 90-105 ° C. In the first embodiment, the pH of the solution in the second leaching stage is again regulated at a value of less than 1, but higher than the value at which arsenic dissolves.

Since acid is consumed in the second leach stage of the process (i.e., when Cu (II) is regenerated), a periodic or continuous supply of an acid such as sulfuric acid, hydrochloric acid, or another acid that does not affect the chemistry may be necessary in the solution in leach stage 16 process. However, the addition of acid depends on whether a sufficient amount of sulfuric acid is formed during the leaching of pyrite. In addition, in the leaching stage 16, the pH is adjusted by adding calcium carbonate to prevent arsenic from dissolving.

At the leaching stage 16, sulfide is oxidized to sulfate and iron is leached into solution in the form of Fe (III) (equation (1)) and typically any gold (or other valuable metal) remaining in pyrite is dissolved. The authors unexpectedly found that the oxidation potential is achievable in a halide solution under the influence of Cu ²⁺ and Fe ³⁺ in the range of 850 mV (compared with SHE) in the presence of oxygen. This potential was sufficient to dissolve gold in the used 8M medium containing Cl ^- , due to the formation of a gold chloride complex.

The resulting pulp containing solid particles from the leaching stage of pyrite 16 is fed to the solid-liquid separation stage 18, in which the solid particles are usually filtered from solution using a known filter unit. The obtained liquid filtrate 20 is fed to the step of extracting the valuable metal 22 “in line”, while the filtered solid particles 24 are removed in the form of tails. To compensate for the loss of water with tails in stage 18, water is added to the top.

The metal extraction step 22 includes one or more columns filled with activated carbon through which the solution flows upward, as in a fluidized bed reactor. Gold (or another valuable metal) solubilized in the solution is adsorbed onto coal, while the liquid discharge stream 26 flows from the column and is recycled to the leaching stage 14. The activated carbon containing gold is then removed or periodically treated and fed as a stream of gold-containing product 28 in the process of extracting gold (for example, by burning a coal product or eluting a column of coal with a cyanide solution).

The liquid discharge stream 26 is recycled to the leaching stage 14 through the iron precipitation stage as the conditioning stage of solution 30. At stage 30, soluble ferric sulfate is precipitated from the oxidation stage of pyrite 16 to remove sulfur and iron from the process by adding limestone and calcium carbonate with the formation of hematite and gypsum (equation (6)). The addition of limestone is adjusted to maintain an iron concentration of about 2 g / l in the solution to prevent the precipitation of divalent copper. Before being removed to the tails, the hematite / gypsum suspension is filtered and the residue washed. After that, the solution is recycled to stage 14.

To prevent the accumulation of contaminants in the whole process, part 32 of stream 26 is recycled as a drain 34 to separate contaminants such as Mn, Cd, Ni, Co, etc. (for example, by controlled precipitation by increasing the pH of the drain solution).

Examples of the first variant of the method

The optimum flow diagram of the first embodiment of the method has been disclosed above, preferred examples will now be described illustrating the first embodiment of the method in accordance with the present invention.

Example 1

For a preliminary assessment, gold extraction from the first concentrate was carried out at three different stages: leaching of arsenopyrite, 1st leaching of pyrite and 2nd leaching of pyrite. The following laboratory test protocols describe the methodology and results of these three steps. CON1 01 refers to the leaching of arsenopyrite and the 1st oxidation of pyrite and CON1 02 refers to the 2nd oxidation of pyrite. The first concentrate was ground to P80 = 30 microns and leached As, followed by oxidation of pyrite.

A task

The challenge was to use IRGP for a single-refractory Au concentrate. An ore sample was delivered to a metallurgical laboratory for grinding and concentration.

Methodology

This experiment was carried out in two parts and was carried out in a 7.5 L tank with titanium insulation. The first part included the leaching of As using a conventional mixer. The second part included pyrite oxidation using a flat-blade turbine and a sprinkler.

Part 1: As Leaching

In a 7.5 L titanium reactor equipped with an impeller stirrer, 3.5 L neutral brine containing 200 g / L NaCl, 50 g / L CaCl ₂ , with a pH <0.5, was prepared. Also prepared were 5 L of an “auxiliary” solution containing 200 g / L NaCl, 50 g / L CaCl ₂ and 75 g / L Cu from CuCl ₂ , with pH <0.5. When necessary, Eh was adjusted between 580 and 600 mV by adding a few grams of copper dendrites. The auxiliary solution was maintained at 80 ° C.

After heating the leach reactor to 105 ° C., an equivalent of 300 g of dry concentrate was added to the brine. When necessary, after 15 minutes, concentrated HCl was added to the suspension to adjust the pH to <0.5, and a sample was taken at t = 0. All acid additives were fixed (time, volume of additive, volume in the leach reactor).

Eh and pH were measured, an auxiliary solution was slowly added at a rate of 2.5 L / h, monitoring Eh so that it did not exceed 530 mV. Every 30 minutes, samples of the solution were taken for analysis on the content of As, Fe, Cu. Eh and pH were monitored every 30 minutes.

When an Eh of 530 mV was reached and it was stable, As leaching was considered complete. The pulp was filtered. The cake was washed twice with hot brine (50 g / L NaCl and pH <1.0), followed by washing with hot water until a clear filtrate was obtained. Kek was dried in an oven all night. Kek was analyzed for the content of S _(total) , S _(ale) , As, Fe, Au and C.

Part 2: Pyrite Oxidation

The 7.5 L reactor was equipped with a flat-blade turbine stirrer and a titanium spray tube. In a leach reactor, 10 liters of concentrated brine were prepared containing 200 g / L NaCl, 50 g / L CaCl ₂ and 75 g / L Cu from CuCl ₂ , with a pH <0.5 and with an addition of 8.8 molar concentrated HCl. The solution was heated to 105 ° C, at t = 0, a sample was taken and the dry cake obtained in the 1st part, where arsenic was leached, was fed into the reactor. After 15 minutes, a sample of the solution was taken to measure Eh and pH. When necessary, technical HCl was added to adjust the pH to <0.5.

Oxygen was supplied at a rate of 2 l / min; Eh and pH were monitored every 30 minutes and samples were taken every hour for analysis on the content of Fe, As, Cu. When the Eh value above 600 mV was stable for 3 hours and the concentration of Fe in the solution did not change, leaching was considered complete. The suspension was filtered. The cake was washed twice with hot brine (50 g / L NaCl and pH <1.0), followed by washing with hot water until a clear filtrate was obtained. Kek was dried in an oven all night. Kek was analyzed for the content of S _(total) , S _(ale) , As, Fe, Au and C.

Example 2

Conducted additional oxidation of pyrite in the residue of the concentrate from example 1.

A task

Analysis of the data and residue from Example 1 showed that the oxidation of pyrite was incomplete when the experiment was completed. In this technique with an improved brine composition, attempts were made to increase the degree of extraction of Au using oxygen to oxidize pyrite.

Result

The second oxidation of pyrite increased the degree of metal recovery, as shown in the following table (based on independent analysis):

Methodology

The 7.5 L reactor was equipped with a flat-blade turbine stirrer and a titanium sprinkler. A 5 L saline solution containing 100 g / L NaCl, 250 g / L CaCl ₂ and 100 g / L Cu from CuCl ₂ , with pH <0.5 and with the addition of concentrated HCl, was prepared in a leach reactor. The solution was heated to 105 ° C, at t = 0, a sample was taken and the dry cake obtained in Example 1 was fed into the As / oxidation pyrite leach reactor. After 15 minutes, a sample of the solution was taken to measure Eh and pH. When necessary, concentrated HCl was added to adjust the pH to <0.5.

Oxygen was supplied at a rate of 2 L / min, Eh and pH were monitored every 30 minutes and samples were taken every hour for analysis on the content of Fe, As, Cu. When the value of Eh above 600 mV was stable for 3 hours and the concentration of Fe in the solution did not change, the flow of oxygen was stopped and Eh was controlled. When Eh remained equal above 600 mV, the oxidation of pyrite was considered complete.

The pulp was filtered. The cake was washed twice with hot brine (50 g / L NaCl and pH <1.0), followed by washing with hot water until a clear filtrate was obtained. Kek was dried in an oven all night. Kek was analyzed for the content of S _(total) , S _(ale) , As, Fe, Au and C.

results

The following results were obtained from the experiments of examples 1 and 2.

Gold recovery results are shown in FIG. 3.

Example 3

In this experiment, the extraction of gold from the second concentrate was achieved in three successive stages: stage 1 — leaching of arsenopyrite and pyrite, stage 2 — leaching of pyrite with oxygen, and stage 3 — leaching of pyrite with chlorine.

A task

Following a search experiment on the leaching of As, in this method the task was to process the second concentrate in an all-in-one process using a brine containing 250 g / l CaCl ₂ and 100 g / l Cu. The solids loading was 200 g / l.

Result

Chlorine oxidation increased Au recovery, as shown in the following table (based on independent analysis):

Methodology

The process was carried out in a 7.5 L reactor equipped with a turbine stirrer. Prepared 5 l of brine having the following composition: 100 g / l NaCl, 250 g / l CaCl ₂ and 100 g / l Cu and the pH was adjusted to <0.5 by adding concentrated HCl.

Part 1: Leaching Arsenopyrite

When the stirrer was rotating at a speed of 90 rpm, its brine was heated to 90 ° C. We took a sample of the solution for future comparison. To the brine was added 1,000 g-equivalent of the dry concentrate "in the form as it was received" (P 80 about 37 microns). A sample was taken at t = 0 and after 15 minutes, Eh and pH were recorded.

Air was supplied to the reactor at a rate of 2 l / min. Eh and pH were monitored every 30 minutes when a sample of the solution was taken for analysis of the content of As, Fe. When the values of Eh and Fe in the solution were stable, the flow of air was stopped. If the Eh value fell by more than 20 mV, the air supply was resumed for 2 hours. When the Eh value did not fall by more than 20 mV, a solid sample of about 100 g was taken and oxygen was supplied instead of air.

Part 2: Pyrite Oxidation

The temperature was increased to 105 ° C. Changed the sample and measurement frequency with an interval of one hour. Under the turbine turbine bottom, oxygen was supplied at a rate of 2 L / min. When Eh and the concentration of Fe in the solution were stable, the oxygen supply was stopped. If the Eh value fell by more than 20 mV, the oxygen supply was resumed for 2 hours. When the value of Eh did not fall by more than 20 mV, the process was considered complete.

The suspension was filtered, the cake was washed twice with acid brine, followed by washing with hot water until a clear filtrate was obtained. The washed cake was dried and weighed. The residue was analyzed for As, Fe, Cu, elemental sulfur, total sulfur, and Au. The last sample of the solution was also analyzed for Au content.

Part 3: Chlorination of Pyrite

To increase the degree of extraction of Au, the experiment was expanded to include chlorination of pyrite using hypochlorite as a source of chlorine. The remainder of the oxidation of pyrite with oxygen was fed into 4 liters of brine having the above composition. The temperature was raised above 100 ° C. and 50 g of hypochlorite was added every 30 minutes. The concentration of Fe was monitored. When the Fe concentration did not increase after the addition of hypochlorite and Eh was stable, the experiment was considered complete.

The suspension was filtered, the cake was washed twice with acid brine, followed by washing with hot water until a clear filtrate was obtained. The washed cake was dried and weighed. The residue was analyzed for As, Fe, Cu, elemental sulfur, total sulfur, and Au. The last sample of the solution was also analyzed for Au content.

results

Gold recovery results are shown in FIG. 4.

The second variant of the method

Before a detailed description of the second variant of the method by example, the second variant of the method will be described with reference to FIG.

In Fig. 5, a concentrate 10 containing valuable metals fed into the process was obtained by mining, grinding and subsequent flotation of sulfide ore. In a second embodiment of the method, the concentrate could be a gold-containing arsenopyrite with a high carbon content (for example, from 2 to 20 wt.% Carbon) or it could have a low carbon content or not contain carbon (for example, a carbon content of less than 2 wt.%). The concentrate was ground in a ball mill 12 and then fed to the stage of oxidation of contaminants in the form of a leaching process of arsenic 14.

A preferred method for leaching arsenic is described in more detail below with reference to FIGS. 6 and 7 and Examples 11 and 12. Leaching can be carried out in one step (for example, if one or more process plants, tanks or reactors are present), but it is usually carried out in as a multi- (two-) stage process. Each stage can have one or more technological units, tanks or reactors operating in direct-flow or counter-current leaching mode, and include (as is known) the use of a drain and a lower product stream.

In any case, a strongly acidic medium is maintained at the leaching stage 14 (the pH is preferably less than 1, since leaching of arsenic from arsenopyrite is favored by a low pH of the solution). An acidic environment can only be achieved by oxidizing the sulfide material (for example, when sulfur in the material is oxidized in solution to sulfate) and / or an impurity-free acid (such as sulfuric or hydrochloric acid) can be added.

In a second embodiment of the method, the Eh of the leach solution is usually maintained at a value of more than 0.4 volts (see FIG. 8) for dissolving a contaminant (e.g., arsenic). As indicated below with reference to Fig.6 and 7 and examples, the leaching process is two-stage. In the first stage of leaching, the Eh of the solution is carefully monitored to facilitate the more preferred oxidation and dissolution of arsenic to its state of oxidation state +3 than to its state of relatively less soluble oxidation state +5, while the pyrite oxidation in the arsenopyrite material essentially does not occur. In the second leaching stage, an oxidizing agent (such as oxygen, air, chlorine, hydrogen peroxide, etc.) is added to increase the oxidation potential of the solution and, as a result, oxidize pyrite (which at the same time oxidizes the remaining arsenic to its oxidation state +5 ) In the second variant of the method and in the second leaching stage, As (V) can either be stored in the solution by the controlled addition of acid (such as sulfuric acid) to reduce the pH of the solution to a value sufficient to dissolve the arsenic, or it can be stored in a precipitated form so so that it is eliminated from the process with pyrite residues.

The process solution is again usually an aqueous chloride solution of divalent copper, preferably having a chloride concentration of 7-8 mol / L. Copper acts again as a leach agent and as a means of electron transfer.

When the sulfide material has a high carbon content (for example, more than 2 wt.%), A masking surfactant can be added to the solution in step 14 to prevent the adsorption of valuable metal leached into the solution by carbon.

At an acid pH of less than 1 and with regulated Eh in the range from 0.4 to 1.25 volts, optimally at about 0.5 volts, arsenic can be oxidized and leached into the solution, preferably in the form of its state with a relatively soluble oxidation state of +3 without oxidizing pyrite sulfide to sulfate, which can adversely affect the properties of the solution.

In the first stage of leaching, the arsenopyrite concentrate is leached for a predetermined period of time (which is indicated in the examples below) until the specified amount of arsenic is leached from the arsenopyrite material (usually about 85% of the total content in the first stage of leaching and an additional 10% of the total content in the second stage of leaching ) In any case, the leached amount is usually determined by the acceptable residual levels in the leached arsenopyrite material, and it is believed that it will be separated and then processed by traditional smelting or calcining methods or removed (as indicated below). The expression “refined arsenopyrite” or “refined sulfide material” should therefore be interpreted in the light of the foregoing.

Thus, in the second variant of the method, the pH and Eh of the solution are adjusted so that arsenic and multivalent varieties of Cu ²⁺ (which oxidize and leach arsenic from the material) remain in the solution and do not precipitate from it in the first leaching stage.

In addition, the process operating conditions are adjusted so that arsenic remains in the solution during the solid / liquid separation step (which separates the refined arsenopyrite solid particles from the solution) until the solution is supplied to the arsenic precipitation stage. 5, this is schematically shown as a thickening stage 16. In the method of FIGS. 6 and 7, said thickening stage is used after the first leaching step. At the thickening stage 16, the solid particles of refined arsenopyrite flocculate (by adding a flocculant to them), while the solid particles precipitate in the form of a lower product stream 18, and the supernatant solution enriched in arsenic and precious metals flows from the thickening stage in the form of a discharge stream 20. In FIG. 5, the bottom product stream or suspension 18 is then fed to a solid / liquid separation step 22, in which solids are usually filtered from solution using a known filter unit.

The resulting liquid filtrate 24 is returned to the discharge stream 20, while the filtered solid particles 26 (i.e. refined arsenopyrite) are fed to the traditional calcination step 28 and the traditional cyanide leach stage 30 to recover the remaining valuable metal in the form of a gold-containing product 32.

Depending on the degree of leaching of the valuable metal in the leaching stage of arsenic 14, gold (and any other valuable metal) can come with a combined liquid stream 34 (combining streams 20 and 24) and can be extracted at the stage of extraction of the valuable metal "on line" 36. The recovery step involves the use of one or more columns filled with activated carbon through which the solution flows upward, as in a fluidized bed reactor. The gold (or other valuable metal) dissolved in the solution is adsorbed by coal, while the arsenic dissolved in the solution passes through the column as a liquid discharge stream 38. The activated carbon containing gold is then periodically removed and fed as a stream of gold containing product 40 (together with a stream of gold containing product) 32) in the process of extracting gold (for example, by burning or eluting a coal product).

The solution 38 (including dissolved arsenic) from the metal extraction stage 36 enters the precipitation stage of contaminants in the form of the precipitation of arsenic 42. Stage 42 is usually carried out at a pH of 1.5-3. At step 42, an oxidizing agent (e.g., air, oxygen, chlorine, etc.) is introduced (e.g., sprayed) into the solution to increase the oxidation potential (Eh) of the solution, which causes the dissolved arsenic to form a precipitate, usually an insoluble precipitate of ferric arsenate (i.e. have FeAsO ₄ or beetroot). When the contaminant includes, for example, antimony, the contaminant may take the form of insoluble ferric antimonate. When a contaminant precipitate forms, an acid is usually generated and, therefore, an alkali may be added to consume the acid and maintain an optimal pH and Eh of the solution. Typically, the alkali is an alkali metal salt such as calcium carbonate, calcium oxide, etc., which has the additional advantage of sulphate separation in the process liquor.

In a second embodiment of the method, the addition of an oxidizing agent and an alkali is controlled to maintain optimal pH and Eh levels at the stage of deposition of contaminant 42 until complete precipitation of all contaminant. After this, the pH and Eh levels of the solution must be restored to those values that are present at the leaching stage 14, so that after the deposition of contaminants, the solution can be recycled to the leaching stage.

At the deposition stage of the second variant of the method, the oxidizing agent causes the oxidation of monovalent copper to divalent copper, as a result of which these varieties are regenerated and their recycling and reuse is possible. Accordingly, the regulation of Eh and the pH of the solution favorably accelerates the re-oxidation of multivalent species while maintaining them in solution at all times so that during the whole process the oxidation state of copper changes from +1 to +2 and it functions as an electron transfer agent and took part in leaching. The regeneration of multivalent species increases the efficiency of the method, simplifies the method and provides a complete material balance of the process.

After arsenic precipitation, the arsenic precipitate is separated from the process liquor in the solid / liquid separation step. 5, this is depicted as an additional condensation step 44, where a lower product stream 46 is formed, consisting of solid particles (arsenic precipitate), which is then fed to the additional solid / liquid separation stage 48. The supernatant discharge stream 50 is discharged from the top of the thickener, used at the thickening stage 44. At the solid / liquid separation stage 48, the arsenic precipitate is usually filtered off using a filter unit to obtain an arsenic waste product 52. The liquid filtrate is returned to the stream Lib 50 as liquid stream 54. The combined liquid stream 56 is then fed to a further step of extracting precious metal 58, e.g., in a column with activated coal, etc. to recover the metal not recovered in step 36. Alternatively, instead of step 36, step 58 may be used. The resulting valuable metal and activated carbon stream 60 is combined with other precious metal recovered streams 40 and 32, while the effluent 62 is recycled to the arsenic leach stage 14, closing the contour of the extraction process.

To process any accumulated contaminant in the process, a portion of the recycle stream 62 can be recycled to the drain circuit 64 to separate the contaminant not recovered in the arsenic deposition step and optionally other contaminants such as Mn, Cd, Ni, Co, etc. .

Now, after the description of the second variant of the method described in detail, the flowcharts of the preferred second variant of the method will be described with reference to Fig.6 and 7.

6 shows a flow chart of a method for processing a single-refractory sulfide material. 6, similar digital references are used to indicate process steps similar or similar to those of FIG. 5. As in the method of FIG. 5, a gold-containing arsenopyrite concentrate 10, 12 is obtained and ground, having a low carbon content or not containing it (i.e., single-refractory). The ground concentrate is then fed to a preferred leaching process. The preferred leaching process has two stages, namely: the first stage of leaching of arsenopyrite (FeAsS) 70 and the second stage of leaching of pyrite (FeS ₂ ) 72.

Arsenopyrite concentrate is fed to the first leaching stage 70, where the leaching conditions are controlled so that only arsenic contained in the concentrate is oxidized and leached into the solution, and not the pyrite component of the concentrate. In this case, the leaching conditions in the first leaching stage 70 are controlled so that the oxidation potential Eh is about 0.5 volts, the pH of the solution is less than 1.0, and the temperature of the solution is maintained at about 105 ° C (although the process can range from 80 up to 105 ° C). The authors found that these conditions are optimal for leaching arsenic into solution. As described in Example 11 below, after about 6 hours of leaching, approximately 85% of the total arsenic content in the arsenopyrite concentrate is leached into the solution.

When a predetermined amount of arsenic is leached into the solution, the solution and refined arsenopyrite containing arsenic are fed to the thickening stage 16 in a manner similar to that described for FIG. 5. The solid particles of refined arsenopyrite flocculate and precipitate in the form of a lower stream of product 18, while the solution of the supernatant enriched in arsenic flows out of the thickener in the form of a discharge stream 20 to the deposition stage of arsenic 42.

In a second embodiment of the method, a solid stream of refined arsenopyrite 18 enters the second leaching stage 72 to leach pyrite. The oxidation conditions in the second leaching stage are more stringent than in the first leaching stage. In this case, an oxidizing agent, such as oxygen, is sprayed into the solution so that the oxidation potential Eh rises to at least 0.6 volts and usually to more than 0.8 volts. In addition, the temperature of the solution in the second leaching stage is maintained at approximately 105 ° C. The pH of the solution in the second leaching stage is still maintained at a value of less than 1.

Since acid is consumed in the second stage of the leaching process (since Cu (II) and Fe (III) are reduced to Cu (I) and Fe (II), respectively), periodic or continuous supply of acid to a solution, such as sulfuric acid, hydrochloric acid, may be necessary acid or other acid that does not affect the chemistry of the process. However, the need to add acid depends on whether a sufficient amount of sulfuric acid is obtained by leaching pyrite. Maintaining a low pH in the second leaching stage also promotes the dissolution of As (V) when necessary.

In the second leaching stage, the sulfide material is oxidized to sulfate, iron is leached into the solution in the form of Fe (III), and part of the arsenic remaining in the arsenopyrite ore is also leached into the solution. The authors noted that an additional 10% of arsenic from its total content can be leached into the solution so that the final residual content of arsenic as a result of the leaching process is 5% or less of its total content in the initial concentrate. This residual content equates to arsenic levels that are low enough for residues safely removed from the process.

The liquor from the second leach stage 72 is supplied as stream 74 to a solid / liquid separation stage 22 similar to that of FIG. 5, in which the remaining solid particles are filtered from the solution and the liquid filtrate stream 24 is returned and combined with the drain stream 20 entering the stage arsenic deposition 42. The solid residues are filtered off in step 22 and then fed as stream 76 to the tails or as filtered solid particles or as a suspension. Alternatively, solid particles can be further processed to recover the remaining metals. To maintain water levels in the process and / or to add to the top of the water lost with the remainder of the process, water may be added in step 22.

In a single-refractory arsenopyrite material and a pyrite material, gold or another valuable metal is not significantly bound to carbon and therefore is usually leached into the solution at both the first and second stages of leaching and, therefore, can be recovered in the process loop.

In the method of FIG. 6, in the precipitation stage of arsenic 42, the pH of the solution is adjusted to about 2-3 (for example, by adding calcium carbonate), and an oxidizing agent, such as air or oxygen, is added to the solution to oxidize the arsenic and transfer it from a soluble state with an oxidation state +3 to its insoluble state with an oxidation state of +5. Since Fe (III) is present in the solution from pyrite oxidation at the second leaching stage, arsenic precipitates favorably in the form of scorodite (FeAsO ₄ ). As an additional advantage, since the sulfide is oxidized to sulfate in the second leaching step, the addition of calcium carbonate can be used to increase the pH of the solution during the precipitation of arsenic and precipitation of sulfate as calcium sulfate.

The arsenic / particulate precipitate, together with the process solution, is then supplied as stream 78 to the solid-liquid separation step 48, in which the solid particles are filtered from the solution. The solids stream of residue 80 typically contains FeAsO ₄ , Fe ₂ O ₃ and CaSO ₄ in a form suitable for disposal (e.g., as a blade). The solids can be removed in suspension and, therefore, water can be added to the top of the process at step 48. Thus, arsenic, iron and sulfur can be successfully recovered in easily removable forms from the initial arsenopyrite concentrate.

The conditions in the arsenic deposition step do not affect the valuable metal leached into the solution in the leaching step, and therefore, the separated solution 56 can now be fed to the valuable metal recovery step 58 in a manner similar to that of FIG. 5. Stage 58 includes one or more columns containing activated carbon onto which valuable metals are adsorbed, typically gold, and the gold-containing product stream 60 is periodically removed from stage 58 and fed to gold recovery (by burning or eluting coal with gold adsorbed on it).

As in the method of FIG. 5, the drain solution 62 from step 58 is recycled to the leach process and a portion of the recycled stream can be removed to the drain circuit 64 to separate contaminants that may accumulate in the process, resulting in a by-product contaminated by-product stream 82.

In a second embodiment of the method, the recycle solution 62 is split to obtain a recycle component 84 fed to the first leach stage and a recycle component 86 fed to the second leach stage. Copper contained in the technological liquor of copper chloride in a state with a oxidation state of +2 is recycled to each of the leaching stages so that it takes part in the leaching of arsenopyrite in the first leaching stage and in the leaching of pyrite in the second leaching stage.

As follows from Fig. 7, it shows a flow chart of a second embodiment of a method for processing a double-refractory sulfide material. 7, similar digital references are used to indicate process steps similar or similar to those of FIGS. 5 and 6. In addition, the upper half (that is, above the dashed line) of the process flow diagram of FIG. 7 is substantially similar to that of FIG. 6 and therefore such process steps will not be disclosed again.

In bivalent arsenopyrite, a valuable metal (for example, gold) is usually associated with carbon and, therefore, such associated gold is not easily leached into the solution at the first or second leaching stages. Accordingly, solids stream 76 includes solids together with an associated carbon / gold component. However, since arsenic, iron, sulfur, and other contaminants are substantially removed during the leaching process to acceptable low levels, the remainder of the solids from the leaching process is now substantially suitable for calcination or smelting in the calcining step 28.

At the calcination stage 28, air and fuel are supplied together with solid particles of material 76, which are calcined in the usual way to obtain a product stream 90, which is then fed to the gold leaching stage 30 in a known manner. Gold leaching is usually carried out by oxidizing the calcined solids with chlorine gas or cyanide (although chlorine gas is preferred because it is less toxic than cyanide). In the second embodiment of the method, part 92 of the solution recycled from the precipitation stage of arsenic 42 is fed to the gold leaching stage 30 to facilitate the overall efficiency of the process.

The off-gas stream 94 from firing stage 28 (typically containing carbon dioxide, sulfur dioxide and other SO _x gases) is supplied to a first gas purification step 96. The first gas purification step typically includes one or more scrubbers in which water and optionally recycled wash water are provided into contact with the gas stream 94. Preferably, the dust present in the gas stream 94 can be filtered. Such dust may include gold chloride (AuCl ₃ ) and trivalent arsenic oxide (As ₂ O ₃ ). Said dust together with other particles can be supplied in the form of solid particles or in solution in the form of stream 98 back to the deposition stage of arsenic 42 for additional extraction of arsenic and gold.

The remaining gases from the first gas purification step 96 are supplied as stream 100 to the second gas purification step 102, typically including scrubbers in which the calcium carbonate in solution is brought into contact with gases containing SO _x . Product stream 104 from the second gas purification step 102 therefore typically contains calcium sulfate and calcium sulfite.

The leached liquor liquor stream 106 containing dissolved gold from the gold leach stage 30 is now fed to a solid-liquid separation step 108 to separate the gold-containing solution from the residue containing solid particles. A stream of solids-containing residue 110 is fed to the tailings for removal, while the gold-enriched solution enters the gold recovery step 114, typically including a column containing activated carbon. Coal and adsorbed gold are periodically removed as stream 116 to recover gold, while the gold-depleted solution 118 is recycled to the arsenic leach / removal circuit to combine with stream 34.

Examples of the second variant of the method

The optimal technological schemes of the second variant of the method have just been disclosed, now preferred examples of the second variant of the method will be disclosed. The following examples processed highly refractory arsenopyrite concentrate from the Bakirchik deposit in Kazakhstan. The task was to develop a method that provided processing of all samples of arsenopyrite ore delivered from the Bakirchik deposit.

Example 4

Concentrate Properties

Methodology:

6 kg of the concentrate was subjected to ultrafine re-grinding. The concentrate had a particle size of P ₁₀₀ 20 microns.

A concentrate with a particle size of P ₁₀₀ 20 microns was obtained in three briquettes, the moisture content of each briquette was determined, and the average value was used as the moisture content of the concentrate.

Briquette 1

Wet sample + paper: 113.84 g

Dry sample + paper: 85.68 g

Paper: 4.83 g

Dry sample: 80.85 g

Moisture Content,%: 25.8%

Briquette 2

Wet sample + paper: 88.35 g

Dry sample + paper: 66.65 g

Paper: 4.83 g

Dry sample: 62.02 g

Moisture Content,%: 25.9%

Briquette 3

Wet sample + paper: 86.41 g

Dry sample + paper: 68.79 g

Paper: 4.85 g

Dry sample: 63.94 g

Moisture Content,%: 21.6%

The determined average moisture content was 24.4%. Based on this indicator, it was calculated that 100 g of dry concentrate was transferred to a sample of wet concentrate weighing 132.3 g.

Example 5

Oxidative Leaching

Tests were carried out on a re-ground sample of P ₁₀₀ = 20 microns to provide initial evidence of leaching of arsenic through an oxidation process. It should be noted that the ore concentrate of the Bakirchik deposit contained arsenic in the form of arsenopyrite. The reaction was designed to determine if arsenic could become soluble (and therefore be selectively removed) when using divalent copper chloride as an oxidizing agent.

Methodology:

Prepared 1 l of a solution containing 80 g / l Cu ²⁺ (in the form of CuCl ₂ 205,13 g), 100 g / l CaCl ₂ , 200 g / l NaCl and 30 g / l NaBr. To the resulting solution was added 140 g of wet concentrate (moisture content ~ 24%, therefore, 105.8 g of dry concentrate) and the resulting suspension was stirred at 105 ° C. For four hours, pH, Eh, and Fe and Cu were measured.

Then the solid particles were filtered using a Buchner funnel and the filtrate was stored for further analysis. The solid cake was washed with a low pH brine (~ 0.5 L, 280 g / l, pH 0.3), the resulting wet cake was weighed, dried in an oven and weighed again. Dry solids were stored for further analysis.

Results and its discussion

The pH, Eh, and total Fe and Cu values recorded over time are shown in the table below.

Analysis of the residue showed an arsenic concentration of 0.66%. It was assumed that the calculated mass loss was 6.5%, while the degree of leaching of arsenic was 82.3%.

The reaction was clearly quick. During the first hour of the reaction, a significant decrease in Eh and pH was observed. After this, the reaction stabilized and no longer progressed.

Example 6

Oxidative Leaching

The objective of this example was to investigate whether fresh liquor would facilitate the further leaching of iron / arsenic from pre-leached material. It was assumed that the processing of solids obtained from the previous leaching will remove most of the arsenopyrite. A fresh stock liquor solution was prepared and the previous leach was repeated using the leached material as a solid charge.

Methodology:

A 500 ml solution was prepared containing 80 g / L Cu ²⁺ (as CuCl ₂ 102.55 g), 100 g / L CaCl ₂ , 200 g / L NaCl and 30 g / L NaBr. To this solution was added 30 g of leached concentrate obtained from the previous oxidation, and the resulting suspension was stirred at 105 ° C. Within four hours, pH, Eh was measured. Then the solid particles were filtered using a Buchner funnel and the filtrate was stored for further analysis. The solid cake was washed with a low pH brine (~ 0.5 L, 280 g / l, pH 0.3), the resulting wet cake was weighed, dried in an oven and weighed again. Dry solids were stored for further analysis.

Samples taken from the solid particles obtained above and from the previous reaction, as well as the starting concentrate, were autoclaved using a mixture of aqua regia / perchloric acid. These solutions were then analyzed for arsenic content using ICP.

Results and its discussion

The pH, Eh values recorded over time are shown in the table below.

Wet cake + paper + filter paper: 72.5 g

Dry cake + paper + filter paper: 40.24 g

Paper + filter paper: 11.5 g

Received dry cake: 28.74 g

The results of the ICP analysis of the arsenic content in the recovered solid particles are shown in the table below.

As in the previous leaching, the reaction proceeded quickly and stabilized within an hour. Again, there was a significant decrease in Eh and pH, together with a decrease in the mass of the recovered solid particles compared to the mass of the solid particles fed into the solution. Based on this fact, it was assumed that the material remaining in the residue from the first leaching was still recoverable. Analysis of the arsenic content both in the initial charge and in the solid residue from both leaches showed a progressive decrease in the arsenic content in the recovered solid particles.

Example 7

Oxidative Leaching

The objective of this example was to improve the conditions used in the leaching stage of arsenic from the ore of the Bakirchik deposit. Having reached the degree of leaching of arsenic contained in the ore from the Bakirchik deposit equal to ~ 65%, the method was improved in order to obtain greater leaching efficiency. The method focused on two directions: firstly, the liquor from leaching was simplified and, secondly, the reaction was carried out at different temperatures and initial pH values to determine the effect of such changes on the increase in leaching efficiency.

Methodology:

A 5 L solution was prepared containing 80 g / L Cu ²⁺ (as CuCl ₂ 1025.64 g), 150 g / L CaCl ₂ (750 g) and 150 g / L NaCl (750 g), and heated to 80 ° FROM. This solution was then divided into three 1.5 L of the solution, with each solution used for oxidative leaching under various conditions of 142.86 g of wet concentrate (moisture content ~ 24%, therefore, 108 g of dry concentrate).

Solution for 1st oxidative leaching: leaching was carried out at 80 ° C.

Solution for 2nd oxidative leaching: leaching was carried out at 100 ° C.

Solution for 3rd oxidative leaching: leaching was carried out at 80 ° C. Initial pH <0.4, Eh> 550 mV.

The pH and Eh of these solutions were measured for 2.5 hours. Samples were taken from each solution at time intervals and analyzed for their iron and copper contents.

The solids were then filtered using a Buchner funnel, and the filtrate was stored for further analysis. The solid cake was washed with a low pH brine (~ 1 L, 280 g / L, pH 0.3) and the resulting wet cake was weighed, dried in an oven and weighed again. Dry solids were stored for further analysis.

Samples taken from the solid particles obtained in each reaction, as well as the starting concentrate, were autoclaved and analyzed for arsenic, copper, and iron using ICP.

Results and its discussion:

Tests of the 1st oxidative leach solution (80 ° C)

Wet cake + paper + filter paper: 173.24 g

Dry cake + paper + filter paper: 105.48 g

Paper + filter paper: 11.5 g

Received dry cake: 93.98 g

Tests of the 2nd oxidative leach solution (100 ° C)

Wet cake + paper + filter paper: 170.3 g

Dry cake + paper + filter paper: 113.32 g

Paper + filter paper: 11.5 g

Received dry cake: 101.82 g

Tests of the 3rd oxidative leach solution (80 ° C, low pH)

Wet cake + paper + filter paper: 172.4 g

Dry cake + paper + filter paper: 108.46 g

Paper + filter paper: 11.5 g

Received dry cake: 96.96 g

The results of the IPC analysis for the content of arsenic, copper and iron in the recovered solid particles are shown in the table below.

These results clearly show that the reaction rate is significantly higher at 100 ° C than at 80 ° C.

Example 8

Iron / arsenic removal process

Methodology:

The liquor obtained from the previous oxidation reaction (10 L) was returned to the vat and heated to 80 ° C with gentle stirring. Upon reaching the indicated temperature, the pH and Eh of the liquor were measured and a sample was taken. Then the liquor was aerated (100 l / h) with stirring, the pH and Eh of the liquor were measured, and then a sample was taken every thirty minutes. After four hours, the removal process was considered complete, the liquor was filtered using a Buchner funnel, and the recovered sediment was removed as a filter cake. Wet cake was weighed and then dried in an oven for twenty-four hours. Then the dry cake was weighed and the sample was autoclaved for analysis.

Results and its discussion:

The pH and Eh values and the composition of the liquor over time are shown in the table below.

Wet cake + paper + filter paper: 257.2 g

Dry cake + paper + filter paper: 128.94 g

Paper + filter paper: 11.5 g

Received dry cake: 117.44 g

Sediment analysis showed the following results:

During the four-hour experiment, essentially 100% of both iron and arsenic precipitated, while the oxidation potential (Eh) was simultaneously restored to a level high enough for use in further leaching. The Fe / As molar ratio was 2.4, while it was expected that FeAsS precipitated along with other Fe-based compounds.

Example 9

Iron / arsenic removal process

Methodology:

The resulting liquor (10 L) was returned to the vat and heated to 80 ° C with gentle stirring. Upon reaching the indicated temperature, the pH and Eh of the liquor were measured and a sample was taken. Then the liquor was aerated (100 L / hr) with stirring, the pH and Eh of the liquor were measured, and then a sample was taken every thirty minutes. After four hours, the removal process was considered complete, the liquor was filtered using a Buchner funnel, and the recovered sediment was removed as a filter cake. Wet cake was weighed and then dried in an oven for twenty-four hours. Then the dry cake was weighed and the sample was autoclaved for analysis.

Results and its discussion:

The pH and Eh values and the composition of the liquor over time are shown in the table below.

Wet cake + paper + filter paper: 257.2 g

Dry cake + paper + filter paper: 128.94 g

Paper + filter paper: 11.5 g

Received dry cake: 117.44 g

Moistened component: 128.26 g (47.8%)

Example 10

Leach regenerated liquor, which is a suspension of low density

Methodology:

A 90 g sample of a wet concentrate (humidity ~ 24%, therefore, 68 g of a dry concentrate) was added to the liquor obtained by the oxidation reaction (1.5 L) and the resulting suspension was stirred at 100-105 ° C. The pH and Eh of the liquor were monitored and a sample was taken every thirty minutes for four hours. After that, the liquor was filtered using a Buchner funnel and the filter cake was removed, the wet cake was weighed and then dried in an oven for twenty-four hours. Then the dry cake was weighed and the sample was autoclaved for analysis.

Results and its discussion:

The pH and Eh values and the composition of the liquor over time are shown in the table below.

Received dry cake: 58 g

Examples of the first and second leaching stages

Example 11

First stage leaching

In this example, the task was to confirm through modeling a continuous method that the technological conditions used for a batch test would be applicable to an industrial process. This experiment also provides material used to oxidize pyrite at atmospheric pressure.

As shown in FIG. 8, a continuous process under conditions similar to the conditions of the batch processes of Examples 5-7 successively gave a gold recovery of 85%.

Methodology:

7.5 L titanium reactors were used, while the discharge from the first reactor was connected to the discharge of the second and then transferred to a collection tank. During a continuous process, a solution was supplied to the first reactor from a supply tank using a peristaltic pump at a rate of 2 L / h. The rate of addition of solid particles was 144 g / h and was achieved by the periodic addition of 24 g of concentrate (in terms of dry matter) to the first reactor every 10 minutes.

Initially, 30 L of a stock solution was prepared containing 80 g / L Cu ²⁺ , 200 g / L NaCl, 100 g / L CaCl ₂ and having a pH <1. 7.5 l of the initial solution was added to each reactor and maintained at 100 ° С, while 360 g of low-grade gold-containing concentrate (30 g / ton) was added in terms of dry matter with a particle size of P ₈₀ = 30 microns and the resulting suspension was mixed and each Eh, pH, and the contents of As, Fe, and Cu were monitored for 30 minutes. After 3 hours, a 100 ml sample of the suspension was taken and filtered on a Buchner funnel and washed with acidic saline. The solids were then dried and analyzed by ICP for copper, arsenic and iron.

After 3 hours, a continuous process was carried out (as indicated) for another 10 hours, with a 200 ml sample of the suspension taken every 2 hours and filtered as described above. The solids were then dried and analyzed by ICP for arsenic, copper and iron.

The results are shown in the following tables:

Analysis of solid particles in the feedstock and discharge of the 2nd reactor

Example 12

The second stage of leaching

In this experiment, the task was to assess the possibility of oxidizing the pyrite component of the residue from the leaching of As with pure oxygen at atmospheric pressure. In this experiment, 500 g of the residue obtained during the continuous leaching test in Example 11 was used.

Pyrite was sequentially oxidized with oxygen at atmospheric pressure and a temperature of 105 ° C. The final recovery of both As and Fe was more than 95%. The content of S _(elem.) (Elemental sulfur) in the oxidation residue was equivalent to the sulfur content associated with arsenopyrite in the concentrate. The results are presented graphically in FIG. 9.

Methodology:

A 7.5 L titanium reactor with a turbine stirrer for gas dispersion and a corresponding titanium gas nozzle on a large yellow heating plate was prepared. In a 7.5 L titanium reactor, 5 L of a concentrated saline solution was prepared having the following composition: 250 g / L NaCl, 50 g / L CaCl ₂ , 20 g / L Cu (from copper chloride) and adjusted to pH <1.0.

A typical sample of the dry residue from As leaching in Example 11 was sent to an independent laboratory for analysis on the content of elemental S, total sulfur, Au, Fe and As.

The stirrer drive was set to a frequency of 80 Hz, the temperature of the solution was raised to 105 ° C, a sample was taken at t = 0, Eh and pH were monitored, and 500 g of the dry residue from the leaching of arsenic from Example 11 was fed into the solution. After 30 minutes, a sample of the solution was taken for analysis on the content of Fe, As and Cu and controlled Eh and pH.

Oxygen was supplied at a rate of 1 l / min. Every 30 minutes for the first 3 hours and every hour after that, Eh, pH, and the contents of Fe, Cu, and As were monitored. When analysis of soluble Fe showed that the increase in its content had ceased, the experiment was considered complete.

The last sample was taken, the suspension was filtered, the filter cake was washed twice with acid brine, followed by washing with hot water until a clear filtrate was obtained. The washed cake was dried, weighed and analyzed for the contents of As, Fe, Cu, C, elemental S and total S.

Experimental results

The following experimental results were observed.

Solid mass after final leaching

Kek:

Wet cake + paper + filter paper: 744.27 g

Dry cake + paper + filter paper: 490.96 g

Paper + filter paper: 66.09 g

Received dry cake: 424.87 g

Weight reduction: 15%

Analysis of solid oxygen fed to pyrite oxidation

Data in dry matter in wt.% Or ppm

Analysis of solid in the remainder of pyrite oxidation with oxygen

Data in dry matter in wt.% Or ppm

The degree of extraction of Fe and As from the concentrate in the residue from the oxidation of pyrite

The recovery of both As and Fe exceeded 95%, as shown in the following tables.

Pyrite oxidation, starting material in comparison with the remainder:

Equivalent Concentrate

The degree of extraction of Fe and As from the concentrate in the residue from the oxidation of pyrite:

The degree of extraction of elemental sulfur from the pyrite oxidation residue showed that the content of S _(elem.) Is equivalent to the content of S associated with AsFeS or 1.4% in the concentrate.

Now, after illustrating preferred methods, those skilled in the art will recognize that the present invention provides the following advantages:

The method can be used to extract valuable metals from sulfide ores and concentrates, which are either difficult to process or impossible to process using traditional available methods / methods, such as smelting and roasting.

The method can be adapted to a high carbon content in such ores, because it is carried out in solution, and, therefore, protective agents can be used to prevent the adsorption of metal by carbon, which, otherwise, could adversely affect the recovery of valuable metal.

The method can be used to remove contaminants from a wide variety of source ores and concentrates, which immediately after their removal can be processed using traditional smelting / roasting methods.

The method makes it possible to remove arsenic, iron and sulfur in easily removable forms from the initial arsenopyrite concentrate, while an easily processed concentrate remains.

The method provides the ability to extract a wide variety of economically significant metals, in particular valuable metals, using simple methods of non-cyanide leaching and separation and adsorption of activated carbon.

The method can be used to treat contaminated residues with their subsequent removal with reduced negative impact on the environment.

Although the invention has been disclosed with reference to a number of preferred options, it is clear that this invention can be embodied in many other forms.

Claims (22)

1. A method of extracting a valuable metal from a sulfide material, comprising the steps of:
preparing an acidic aqueous halide solution having an oxidizing potential sufficient to oxidize the sulfide material and dissolve the valuable metal in the solution;
adding material to the acidic aqueous halide solution so that the sulfide material is oxidized and the valuable metal is dissolved; and
separating the valuable metal from the oxidized sulfide material, in which the metal halide is selected to function as polyvalent species during the oxidation of the sulfide material. 2. The method according to claim 1, in which the sulfide material is contaminated with arsenic, the acidic aqueous halide solution has a pH at which arsenic precipitates in such a way that when the material is added to the acidic aqueous halide solution, the sulfide material is oxidized, the valuable metal dissolves and precipitates arsenic, after which separates the valuable metal from the oxidized sulfide material and precipitated arsenic. 3. The method according to claim 1 or 2, in which the solution containing the valuable metal is separated from the oxidized sulfide material and precipitated arsenic (when present) in the solid / liquid separation step, and the valuable metal is recovered from the solution in the metal extraction step, in which the valuable metal is adsorbed by activated carbon in one or more carbon-containing columns, and after the adsorption of the valuable metal by activated carbon, the carbon is eluted with a cyanide solution and the eluate is fed to the electrolysis step to recover the valuable metal. 4. The method according to claim 3, in which the stage of metal extraction is carried out on the line after the stage of separation of solid-liquid and before recirculation of the solution to oxidize the sulfide material. 5. The method according to claim 1 or 2, in which the valuable metal to be extracted is gold, silver, platinum or another platinum group metal. 6. The method according to claim 1 or 2, wherein the aqueous halide solution is a soluble metal halide solution having a halide concentration of about 8 mol / L, wherein the halide is a chloride or a mixture of halides containing chloride and bromide. 7. The method according to claim 6, in which the metal in the solution of the dissolved metal halide is copper and / or iron. 8. The method according to claim 1 or 2, in which the stage of oxidation of the sulfide material includes one or more stages of leaching so that:
(i) for a non-impurity single-refractory pyrite material, the stage of oxidation of the sulfide material included one leaching stage in which the pyrite material is oxidized and the valuable metal is dissolved at the same time; or
(ii) for contaminated single-refractory or double-refractory pyrite material, the stage of oxidation of the sulfide material included a two-stage leaching process in which the solution from the first leaching stage is fed to the second leaching stage. 9. The method of claim 8, in which: (ii) the pyrite material is arsenopyrite, and in the first stage of leaching, the oxidation potential is regulated to leach arsenic into the solution and the pH of the solution is adjusted so that immediately after leaching, arsenic precipitates as arsenate ferric iron, and in the second of the leaching stages, the pyrite component is leached and the pH of the solution is adjusted to maintain arsenic as a precipitate of ferric arsenate so that arsenic is removed from the process with oxidized sulfide m material, and
in the first leaching stage, the material is brought into contact with the solution at Eh of about 0.7-0.8 V, sufficient to leach contaminants and dissolve the valuable metal, at a pH of less than 1, but more than about 0.5, in order to precipitate arsenic immediately after its leaching, and at a solution temperature of about 80-105 ° C, and where
in the second leaching stage, the material is brought into contact with a solution having an Eh of about 0.8-0.9 V, sufficient to leach pyrite, the pH of the solution is less than 1, but more than about 0.2, in order to precipitate arsenic immediately after it leaching, and at a temperature of the solution from about 90 to 105 ° C. 10. The method according to claim 3, in which, after the extraction of the valuable metal, the stage of conditioning the solution is used to precipitate ferric sulfate and thereby control the level of these varieties in the process,
at the stage of conditioning the solution, limestone and calcium carbonate are added to the solution to form a hematite / gypsum precipitate, which is then filtered and removed with the residue consisting of solid particles from the leaching stage (s). 11. The method according to claim 1 or 2, in which, if a high level of carbon is present in the sulfide material, a surfactant is added to the solution during the oxidation of the sulfide material to prevent the adsorption of the valuable metal on the carbon contained in the material or in the solution activated carbon is added during the oxidation step of the sulfide material to preferentially adsorb the valuable metal with activated carbon, wherein
the surfactant is one or more organic solvents including kerosene or phenolic ether. 12. The method of removing contaminants from contaminated with impurities sulfide material, comprising the steps of:
mixing the material in an aqueous solution in which polyvalent species with a relatively high degree of oxidation oxidize the contaminant, making it soluble in the solution, and a refined material containing contaminants is obtained and reduced to a relatively lower oxidation state; and
removing contaminants from the solution while regenerating the polyvalent species to their state with a relatively high oxidation state, the contaminant being removed from the solution by precipitation at a separate precipitation stage by introducing an oxidizing agent into the solution,
The pH of the solution in the precipitation step is usually maintained at about pH 1.5-3.0. 13. The method according to item 12, in which the multivalent species are dissolved copper and / or iron. 14. The method according to item 12, in which the pollutant is oxidized and leached into the solution in a two-stage leaching process, where in the first stage of leaching, the oxidation potential is regulated to oxidize essentially only the pollutant to make it soluble in the solution, and to the second the leaching stage, the oxidation potential is increased for the oxidation of sulfide in the contaminant-containing refined material, and the contaminant-containing refined material is separated from the solution after the first tadii leaching and fed to the second leaching step, and the solution was separated from the contaminants-containing material is refined after each leaching step to remove it from contaminant deposition. 15. The method according to 14, in which the sulfide material is a pyrite material, and in the first stage of leaching, the pollutant is oxidized in an acidic aqueous solution with a pH of less than 1 at an Eh of a solution sufficient to oxidize the pollutant in the solution, but not for leaching, essentially pyrite, and at temperatures up to about 105 ° C and in the second stage of leaching, the pyrite material is also oxidized in an acidic aqueous solution with a pH of less than 1, but at a higher solution Eh, sufficient to leach pyrite, and at temperatures up to about 105 ° S When in use,
In the second leaching step, an oxidizing agent such as oxygen, air, chlorine gas, hydrogen peroxide is added to the solution. 16. The method according to item 12 or 13, in which the solution is recycled to the process and it is a solution of dissolved metal chloride having a chloride concentration of about 8 mol / l, and the metal in the solution of dissolved metal chloride functions in the form of polyvalent species. 17. The method according to item 12, in which the sulfide material has a high level of carbon present in it, and a surfactant is added to the solution during the oxidation of the contaminant to prevent the adsorption of valuable metals by the carbon contained in the material. 18. The method according to item 12, further comprising one or more stages of metal extraction to extract the metal leached into the solution with a contaminant, and / or that is present in a refined material containing a residual contaminant, so that when
the sulphide material is a double-ore ore comprising carbon, then after the final leaching step, a metal extraction step is provided to recover the metal present in the refined material containing the residual contaminant that has been adsorbed by the carbon, and
a metal recovery step, which is a traditional firing or smelting process, optionally leaching with chlorine or cyanide after firing or smelting to recover the metal remaining in the calcined solid particles of the material. 19. The method according to p. 18, in which before and / or after the deposition of contaminants provides a stage of metal extraction on the line, designed to remove metal that is leached into the solution at the stage of leaching,
prior to metal recovery, a series of material separation steps are provided for separating the contaminant-containing refined material from the solution, wherein
the solution separated after the leaching stage or after each leaching stage is supplied to the stage of extraction of the contaminant, and the separated refined material is fed to the extraction or removal of metal. 20. The method according to 14, in which after the deposition of the contaminant provides a stage of separation of the contaminant, designed to remove contaminants from the solution before recycling the solution to the leaching stage. 21. A method for removing a contaminant from an impurity-contaminated sulphide material, the method comprising the steps of: mixing the material in an aqueous solution having an oxidizing potential, regulated essentially to oxidize only the contaminant to make it soluble in the solution, whereby a refined material containing contaminants is obtained; and
separating the solution from the contaminant-containing refined material, the contaminant being oxidized and leaching into the solution in a two-stage leaching process, where in the first leaching step the oxidation potential is regulated to oxidize substantially only the contaminant to make it soluble in the solution, and in the second leaching step the oxidizing potential is increased to oxidize the sulfide in the contaminant-containing refined material, and
the contaminant-containing refined material is separated from the solution after the first leaching step and fed to the second leach stage, and the solution is separated from the refined contaminant-containing material after each leaching step to remove contaminant by sedimentation. 22. The method according to item 21, in which the sulfide material is a pyrite material, and in the first stage of leaching, the contaminant is oxidized in an acidic aqueous solution with a pH of less than 1 at an Eh of a solution sufficient to oxidize the contaminant in the solution, but not for leaching, essentially pyrite, and at temperatures up to about 105 ° C and in the second stage of leaching, the pyrite material is also oxidized in an acidic aqueous solution with a pH of less than 1, but at a higher solution Eh, sufficient to leach pyrite, and at temperatures up to about 105 ° S When in use,
In the second leaching step, an oxidizing agent such as oxygen, air, chlorine gas, hydrogen peroxide is added to the solution.

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